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By Steve Hansen

Heap leach mining has long been an economical method for mining oxide copper ore, and with the U.S. Bureau of Mines' discovery of a way to economically strip gold and silver from carbon, it has become the standard in the gold and silver indus­try as well. Sulfide leach in copper recovery has always been a byproduct of oxide leach mining, although the methods and recovery rates have increased with technology. SX-EW has kindled a new interest in large deposits and dormant dumps of low grade sul­fide ore. Gold ore bio-oxidation was developed in the 1980's, but not tried on a large scale until US Gold built the Tonkin Springs Mine. This mine was dedicated to the large scale production of low grade sulfide ore that could be recovered through the newly discovered bio-oxidation method. On a small scale the method worked, but as time went on, problems with the process were costly and recovery was poor. US Gold did not have the finan­cial backing to perfect the process, and the mine was closed in 1990. Since that time, Newmont Gold Company has taken the lead and has dedicated a lot of resources to the perfection and implementation of this method. Some small mines have discov­ered that if they can use this technology they may be able to dou­ble and triple current ore reserves, because they will be able to mine deposits previously considered not to be economical.

Jan 1, 1998

By I. S. Parrish

Reserve audits have increased in recent years as more companies, especially juniors, enter the ranks of producers. Corporate and individual investors are seeking corroboration of stated reserves. The reserve audit satisfies this need. An audit assures investors that the reserves, as presented, are a reasonable and accurate approximation of the quantity and grade of the economically viable material present. The reserve audit verifies that there are no fundamental errors in logic, style or the data base, and that the reserves, as given, may be used for mine planning or economic forecasting. Reserves that pass an audit may be used with confidence in preparing cash flow projections and in deriving net present value or internal rate of return on any given project. This article describes what a consultant does while performing the reserve audit. In performing an audit there are four general areas of concern that must be checked: definitions, parameters, data and procedures. In each of these, the auditor must convince himself that the reserve study has complied with governmental regulations and guidelines, professional standards and industrial practices. Any deviation from these norms must be identified and its effect upon the appropriateness of the reserves evaluated. By concluding that a reserve study passes his audit, the reviewer is stating that there are no fundamental errors in logic or style and that spot checks have revealed no substantive errors in procedure or mechanics. The responsibility for the reserves remains with the entity that compiled the original study.

Jan 1, 1990

By Jim Peregoy

INTRODUCTION Frontier-Kemper Constructors, Inc. is an Evansville, Indiana based heavy construction contractor engaged in projects for both the public and private sectors of the underground construction industry. Construction activities range from conventional shaft, slope, tunnel and chamber excavation through state-of-the-art mechanical excavation, ground freezing, standard support and structural concrete operations. Conventional excavation and concreting equipment, as well as raise boring machines, road headers, tunnel boring machines, and freeze plants are owned, operated, maintained and modified "in-house" to suit the needs of individual projects. The balance of this paper wilt present a general overview of the Frontier-Kemper planning and construction operations on the recently completed Phase II of the U.S. Army Corps of Engineers, John Day Dam, Juvenile Fish Bypass Project. General Project History The U.S. Army Corps of Engineers is responsible for designing, constructing and maintaining various navigation and impoundment structures on the nation's inland waterways. Maintenance involves both the work required to maintain the structure in fully-operable condition and the ongoing work required to maintain a positive impact on the environment. John Day Dam is located on the Columbia River approximately 95 km (60 miles) upstream of The Dalles, Oregon. It was constructed to: provide electric power to the Pacific Northwest; provide Locking for up-and-downstream bound commercial and private watercraft; act as a water impoundment structure to control flooding conditions; and to provide a source of irrigation water

Jan 1, 1987

By Philip A. Mattson

INTRODUCTION Engineers involved with rock mechanics know the importance of discontinuities within a rock mass and their effect on its behavior when subjected to loads from structures or construction excavation activities. Various geophysical tools have been developed to provide comparative information on subsurface conditions, such as electrical resistivity, assorted seismic test methods, acoustic tests, etc. While these investigative means have proved invaluable in numerous situations, especially those where site alternatives are being evaluated, they shed little light on the difficult problem of quantifying the rock strength. For that purpose, a laboratory test program of samples obtained in the field is usually undertaken, often in conjunction with a variety of in-situ tests. While standard samples provide sufficient information on the rock properties, questions on the macro-structure of joints, faults, fissures, and their infillings often go unanswered. Presently, attempts are being made to incorporate into rock sampling technology new techniques which are directed towards providing better information on the incidense and orientation of discontinuities. One such device incorporates a scribe knife within the core barrel which scores the sample lengthwise as it is cored. The orientation of the scribe is known, thereby relating the bedding, planes, fractures, etc. to a spatial framework. A second approach, called the Integral Sampling Method, is the subject of this paper. Stated briefly, this method utilizes a grout injection to bond fractured rock to a steel reinforcing bar at the center of the coring. The bar is intended to act as a

Jan 1, 1979

By R. M. Balderrama

Bench-scale laboratory flotation and gravity separation tests were conducted on tailings samples to remove the sulfidic material, mainly pyrite, the source of acid generation. A combination of xanthate collectors, promoters, and frothers were used during flotation. Pyrite recoveries were in the range of 42 to 97%, depending on the tailings being treated. Sulfide sulfur contents of 12.8 to 0.37% in the feeds were reduced to 8.68 to 0.01% in the treated tails. Flotation tests were conducted to evaluate the ARD potential as a function of percent pyrite remaining in the tails. Standard ARD predictive tests were used to evaluate the success of the treatment. In some cases, removal of only 50 to 55% of the pyritic fraction was required, thereby demonstrating the conversion of the waste material from acidic to non-acidic. A laboratory vanning table was used to evaluate the amenability of sulfide-bearing tailings samples to gravity separation. The pyrite recoveries varied from 38 to 86%. The sulfide sulfur content of 12.8 to 0.37% in the feeds was reduced to 6.0 to 0.14% in the treated tails. Predictive ARD tests performed on tailings before and after the gravity removal of pyrite showed that most samples changed from "probably" acid producing to the "uncertain" range. Data reported and analyzed in this paper demonstrate source control for ARD prevention through the removal of metal sulfides from sulfide-bearing wastes using conventional flotation and gravity separation techniques.

Jan 1, 1995

By Mike Horn, Keith Meade, Marvin L. Smallwood

Technological changes in programmable logic controllers (PLC)'s and the need to improve the architecture of Mine Wide Monitoring Systems (MWMS) has resulted in the design of a PLC based system at the Virginia Division of Island Creek Coal Company. The system which is presently in the installation phase, is a joint effort of Island Creek Coal Company, Westinghouse Electric Supply Company and Mining Controls Incorporated. The system uses an ethernet based network for graphic display of system data operating under the Genesis software package. Telemetry for the system utilizes the Seimens/Westinghouse L1 network operating over two twisted pairs. Remote telemetry uses fiber optic links and modems operating over standard telephone lines. 500 and 2000 series PLC's provide distributed data processing for the system with hierarchial processing operating under the Genesis software on the ethernet network for individual mine site monitoring along with the remotely located division office monitoring. System control and data acquisition is provided for underground MWMS operations and on the surface for a vast methane gathering system that is used for degasification of the mines and collection of pure methane for sale to the Columbia Gas Company as natural gas. The inability of industrial MWMS companies to keep up with the vast changes in technology, along with economic considerations associated with expanding the existing MWMS created the need for the development of the PLC based system to replace the older industrial monitoring systems.

Jan 1, 1993

By J. W. Langhans, B. D. Blake

The U. S. Bureau of Mines has been investigating the application of thiosulfate to recover gold directly from low grade refractory carbonaceous ores for the past four years. Statistical experimental methods were used to determine the effects of various leaching parameters on both gold extraction and thiosulfate consumption. Seven variables were investigated, including thiosulfate, copper, ammonia, sulfite, and sulfate concentrations, leaching time, and leaching atmosphere. This information was used in face-centered cubic (FCC) surface response experiments to define the optimum leaching conditions for a low grade ore. Two models were generated from the data to predict gold extraction and thiosulfate consumption as a function of the leaching parameters. The models predicted a gold extraction of 67 pct and reagent consumption of 2 kg [S2032/mt ore at 0.25 M S2032-, 0.0014 M Cu2+, 0.27 M NH,OH, 0.00625 M SO;-,] and leaching for 16 h. Actual tests resulted in 62 pct gold extraction with 0.93 kg S2032- consumed/mt ore. Subsequent column leaching tests have further demonstrated the viability of applying this technology to recover gold from carbonaceous ores without first pretreating the material. Gold extractions in the column tests ranged from 69 to 78 pct with thiosulfate consumption ranging from 0.4 to 5.4 kg/mt ore. These results compare to <5 pct gold extraction using standard cyanide leaching conditions.

Jan 1, 1996

By S. P. Banerjee

By international standards, Indian coal seams are not very gassy. In some of the major coalfields such as Singareni, Wardha, Pench, Chirimiri, Rewa and Talcher Coalfields practically no methane is emitted from the coal seams. Even in the coalfields of Ranigang, Jharia, Bokaro and Karanpura, all of which are situated in the basin of the Damodar river, only a few mines are faced with excessive gas emission problems. But the number of such mines is increasing as deeper reserves are being worked and already the exploitation of some of the best quality coal reserves in the country is held up pending measures for control of methane emission in these mines. Even in mines having no previous history of gas emission, methane has been met near dykes and faults. This was dramatically revealed by an explosion in Dhori colliery in 1965 wherein 268 persons got killed. Until that time, Dhori colliery was considered as a non-gassy mine but on that fateful day gas came to the workings through a freshly exposed fault plane. Similar incidences of ignition of firedamp in so-called non-gassy mines have also occured in the past, for example, in Chapui Khas and Samla Dalurband collieries, but the casualty figures were not as high. After the Dhori explosion, all Indian coal seams have been declared to be gassy, whether presence of methane is detected in the workings or not.

Jan 1, 1980

By Kim A. Lapakko, William W. III White

The ASTM standard method D 5744-96 was developed to provide a well-defined kinetic test protocol, conducted under controlled reaction conditions, with the intent of generating reproducible results. The US Bureau of Land Management, American Assay Laboratories, and the Minnesota Department of Natural Resources cooperated in a intralaboratory and interlaboratory testing of the protocol. Intralaboratory testing was conducted on duplicate samples of gabbro (8) and siltite-argillite (5) rocks for periods of 20 to 125 weeks. The difference between replica& intralaboratory pH readings on drainage from the gabbro for the drip-trickle leach alternative typically did not exceed 0.1 units, although differences as high as 0.4 units were observed. The difference from the mean for rates of sulfate, calcium and magnesium release was typically within 10%. Intralaboratory results for the siltite-argillite samples were generally similar, although a one unit pH difference was observed for one sample. Three different laboratories tested a gabbro sample for 125 weeks, with a maximum difference from the mean pH of 0.5 units. The percent difference from the mean for sulfate and calcium release rates was typically within lo%, although values as high as 30% were observed. Data on intralaboratory and interlaboratory agreement on the flood rinse alternative of the method were limited. Replication on the sample tested was comparable to of better than that observed for the drip-trickle method testing of the same sample.

Jan 1, 2000

By L. Barry Phelps

The complexity of surface mining operations is increasing as attention is directed towards mining larger tracts of lower grade, higher cost deposits. Large tonnage requirements from these mines necessitate utilization of many large pieces of equipment. This equipment must be utilized efficiently, with production maintained at or near designed capacities. Basically, the problem is to select, size, and schedule equipment to maximize profits and minimize adverse environmental impacts (Anon., 1977). The time, manpower, and cost limitations required to analyze a wide variety of situations and alternatives in planning surface mines have been sidestepped with the application of simulation methods using a digital computer. Since a simulator can be programmed to represent the situations which are to be analyzed, many alternatives can be evaluated in a relatively short time. Input variables are easily changed and the resulting changes which take place in the system can be analyzed. When standard methods of analyses are applied, weeks and many man-hours are involved in analyzing large-scale mining operations. At the same time many simplifying assumptions must be made since these operations are dynamic and transient. Hence, it becomes difficult to make sound engineering decisions and take proper and timely corrective action in the absence of simulation methods. This chapter is concerned with the application of one such method, namely, the open-pit materials handling simulator (OPMHS) developed by the Mining Engineering Section, The Pennsylvania State University (Manula, Albert, and Ramani, 1977).

Jan 1, 1979

By M. P. Hardy, C. J. H. Brest van Kempen, H. N. Maleki

To evaluate the influence of uniformly tensioned roof bolts on roof stability, mechanically anchored roof bolts were installed using the conventional bolter and a specially developed thrust and torque controlled retrofitable box mounted on a bolting machine. Thirteen thousand bolts were installed over 19 crosscuts in a 4 heading development in an operating coal mine. Geological , bolt and mine parameters were recorded at bolt installation, and bolt tension histories, roof sag and other changes evident in the rooms, crosscuts and intersections were recorded over a 2 month period. In all, 24 variables and 11 responses were quantified at 67 bolt clusters. These data were analyzed statistically to determine the significant variables influencing roof stability. As the mine development was safe and stable during the monitoring period, roof sag after two months was used as an index to quantify "stability." The variables of significance were shown to be coefficient of variation of bolt tension, bolt anchorage capacity, roof lithology, roof shape, standard deviation of bolt tension, and dip of roof lithologies. It was concluded that of the variables with which the mine operator has control, bolt tension uniformity was the most significant parameter at this mine. Other variables, such as intersection type and time between mining and bolting , were not significant, whereas, some other potentially significant variables such as horizontal stress or roof joint frequency did not vary sufficiently for their influence on stability to be evaluated.

Jan 1, 1986

By R. Poulin, R. Pakalnis, D. Peterson

Introduction A comprehensive project has been completed at the University of British Columbia that focused on the development of a commercially competitive composite cable bolt. Laboratory evaluation consisted of pull tests, preliminary shear tests, grout column testing and scanning electron microscope evaluation of the composite cable bolt. Field implementation at five underground mines throughout Canada were completed for technical and overall performance evaluation of the new product. The project began in 1989 under the sponsorship of HDRK Ltd., a consortium of mining companies including Inco, Noranda and Falconbridge Ltd. The objective was to develop a cuttable cable bolt that would serve as support for a continuous miner within a hardrock environment (Mah, 1994). The bolt was required to be cuttable by a continuous roadheader without adverse effects to equipment, personnel and the milling process. From the beginning, the focus was to employ existing cable bolt installation technology (Goris et al., 1994) whereby the cable bolts are grouted within a 4.8-cm-diam (1.08-in.diam) or larger employing a portland cement grout having a water: cement ratio between 0.35:1 to 0.4:1. The prototype, developed from Phase I, was technically comparable to that of conventional steel cable bolts. However, high cost and limited accessibility hindered its commercial acceptance. The second phase of the project was sponsored by the British Columbia Science Council and Pacific Pultrusions Ltd. of British Columbia to locally develop a glass fiber cable bolt. This phase of the project has resulted in the DAPPAM cable bolt (Pakalnis et al., 1994). DAPPAM Cable bolts are the second most used support method in Canadian underground mines after mechanical rock bolts. It is projected that 870,000 m (2.9 million ft) of cable bolts are installed annually in Canada (Poulin et al, 1994). The objective of the research program was to develop a glass fiber cable bolt with similar pullout strengths to the conventional steel cable bolts (267 kN - 60,000 lbf). More than 200 laboratory pull tests were conducted employing standard cable bolt test procedures, as outlined by the US Bureau of Mines (Goris, 1994), to assess optimum rod and surface composition, geometry, number required, pullout and shear characteristics, among others. This resulted in the DAPPAM cable bolt (Fig. 1). It consists of 10 individual 0.6cm-diam (0.25-in.-diam) rods enveloping a 2.2-cm (0.8-in.) outside diameter, high-density polyvinyl grout tube. The composite used in the construction of the rods is comprised of about 65% glass fiber within a 35% polyester resin matrix (by volume). Through pultrusion (a continuous fabrication technique), the fibers and resin are combined into a composite material. The strength of the overall composite is largely governed by the material properties, the alignment [ ] of the fibers relative to the applied load and the density and interaction of the overall components within the composite. A further modification to the composite was the addition of a sand grit to the surface of the individual rod. This increased the overall bond strength between the rod and the cement-based grout, used to solidify the cable bolt to the rock mass. The grout tube has an inside diameter of 1.9 cm (0.75 in.) and is rated at a pressure of 1.72 MPa (250 psi). During installation, the grout is pumped through the annulus of the grout tube thereby filling the hole from toe to collar. Laboratory and field trials concluded that a minimum drill hole diameter of 4.8 cm (1.08 in.) be used to ensure optimum grout coverage. Table 1 compares the design strengths (as tested) for DAPPAM vs. the ultimate strengths for a conventional 1.6cm (0.6-in.) steel cable bolt. The characteristics for the steel cable have been derived directly or indirectly by Goris et al. (1994). A design tensile strength of 289 kN (65,000 lbf) was proposed for the composite developed. However, recent tests have indicated maximum breaking strengths of more than 400 kN (90,000 lbf) (Fig. 2). Figure 2 shows the load vs. displacement pull out curves for steel and DAPPAM cable bolts. The methods of testing were standardized so as to be able to compare the results directly. The shear strength of the fiberglass cable is largely a function of the surface area sheared. Shear strengths of 89 kN (20,045 lbf) were obtained for the configuration shown in Fig. 1 when sheared normal to the axis of the cable. The testing procedure was similar to the shear tests performed on steel by Goris (1993). It is important to reinforce the implication that critical bond strength has on the success of an overall support system.

Jan 1, 1995

The public relations problems of users of explosives have greatly increased within the past few years. One reason is that explosives and blasting agents are being consumed in increasing quantities. The principal reason, however, is that more people have become physically affected by blasting. Mines, quarries, and construction sites that were formerly in relatively remote locations now find themselves near congested areas. New express traffic ways being constructed to connect the suburbs with downtown areas are being built through the middle of long-established business and residential neighborhoods. The public becomes aware of blasting by two physical sensations-vibration and noise. Vibrations are usually so slight as to be negligible, and in most cases the only sensation noted is noise. The reduction of vibrations to this low level has been primarily achieved by the use of blasting practices and explosive products recommended by vibration engineers based on seismograph recordings. Unfortunately, very little scientific attention was given to the accompanying physical sensation-noise.

Jan 1, 1962

By Wade E. Cooper

INTRODUCTION How accurate are past and present employee radondaughter exposure records of underground uranium miners employed in the United States? This often-debated question is essential for future substantiation of safe exposure limits. An apparent discrepancy between company-reported exposures and Mining Enforcement and Safety Administration (MESA) projected exposures was detected in 1977. For these reasons a need for an updated comparison of these exposure data was indicated. This paper gives some of the conclusions of the earlier study and compares more recent exposure records compiled by the Atomic Industrial Forum, Inc., with projected exposures based on sampling by Federal mine inspectors. EARLIER STUDY In its 1977 Annual Report (U.S. Department of the Interior, 1978), MSHA&apos;s predecessor, the Mining Enforcement and Safety Administration (MESA), reported that there was "an apparent discrepancy between Federal inspection results and company records." Both company records and MESA&apos;s projections from samples taken during routine Federal inspections indicated reductions in the average exposure of underground uranium miners from 1975 to 1977, but the MESA projections were over 4 times higher than the company-reported averages. This apparent discrepancy however, was based on a comparison of exposure data reported for all U.S. underground uranium miners. This projection more closely represented the average exposure of U.S. underground uranium mine production workers who worked 1,500 hours or more during the year. Exposures of such workers are reported each year by the Atomic Industrial Forum, Inc. (AIF) in summaries of exposure data reported to the AIF by uranium mining companies throughout the United States. (The AIF exposure summary for 1979 appears as tables A-1 and A-2 in the appendix of this paper.) Assuming that the average exposure for each exposure range category is the midpoint of each exposure range category, table 1 compares the estimated average exposures for U.S. underground uranium mine production workers who worked underground 1,500 hours or more each year in 1975 through 1977 with the exposures projected by MESA for those years. [Table 1. - Average Exposure and Projected Average Exposure for U.S. Underground Mine Production Workers Who Worked Underground 1,500 Hours or More During the Year. Company, MESA?&apos; Reported- Projected Year (WLM) (WLM) 1975 1.59 5.68 1976 1.84 4.64 1977 1.68 4.08 1 Atomic Industrial Forum, 1976, 1977, 1978. 2 U.S. Dept. of the Interior, 1978.] Table 1 indicates that, even after adjustment to ensure better comparability an apparent discrepancy between Federal inspection results and company reported exposures for 1975-1977 exists; however, the apparent discrepancy diminished over the 3 years. Slade, 1977, explained some of the discrepancy between company records and MESA projections of miners&apos; average radon-daughter exposures as follows: 1) Concentrations of radon daughters in some work areas can vary greatly during any one day. A variation from 0.3 WL to 17.0 WL has been measured in the same stope on the same day. 2) Seemingly simple abatement problems indicated by the regular Federal and State inspections were solved simply by manipulating the mine ventilation. 3) The methods used by mine operators to compute cumulative exposures were such that high radiation readings were seldom or never reflected in the records. For example, a work area sampled on Monday indicated a radon-daughter level equal to 0.2 WL and this was recorded. It was sampled again on Wednesday of the following week and the level was 2.2 WL. The miners were withdrawn or told to fix the ventilation, and when this was accomplished the area was sampled and found to be at 0.2 WL again. Although the miners could have been working in the higher concentration up to 6 days, this reading might never be reflected in their records. If it was recorded, only a fraction of the day on which it was discovered would be entered into the cumulative exposure calculation (time-weighted average). 4) Some of the mines visited used a mine average radiation concentration, and every employee working underground was given the same exposure per unit of time spent underground. As a result of the 1977 study, more stringent sampling and recordkeeping standards were proposed and public hearings held in 1977. The resulting new and revised health standards on radon-daughter sampling and exposure recordkeeping became effective August 30, 1979 (Mine Safety and Health Administration, 1979). Prior to these new regulations, radondaughter sampling requirements were on an "as often as necessary" basis (Code of Federal Regulations, 1978). The new regulations required practically all active work areas in underground mines to be sampled at least once every 2 weeks, with many areas requiring weekly sampling. They also required calendar-year exposure records of all underground

Jan 1, 1981

By Stephen J. Swan

Feed-forward control (i.e. the forward flow of information in a control system ) is a simple and flexible approach to over-all control of a ball mill-rake classifier grinding circuit. The system described here uses mill power, feed rate, mill dilution water flow rate, and classifier dilution water flow rate only, to control mill load, throughput, mill solids, and classifier solids as well as classifier overflow grind. The mill load is controlled by a simple feed-back control loop, while mill solids and overflow grind are controlled within close limits by a unique non-linear feed-forward water control system. In effect the control follows a limited mathematical model of the grinding circuit. The configuration minimizes deviations in overflow grind due to changing ore conditions, but since it is open-loop, and the model is not complete, it is affected by changes in total ball charge and liner wear. These effects are fortunately gradual, and changes in the grind can usually be kept to within ± 1% of a target continuously for several days without adjustment. Control adjustments are made with feed-back from laboratory composites. This control system has won enthusiastic operator acceptance; it is easily started up and shut down; grind-outs have been eliminated along with flotation machine flooding and mill spills. Standard measurements and components are used throughout providing a trouble-free system design at a moderate cost.

Jan 1, 1970

By E. Maré

Some of the members will probably ask themselves what close relationship exists between hydro-metallurgy and a wet scrubber, but we are sure that as soon as deeper thought is given to the matter it will be realised how vital scrubbing, operations are in metallurgical plants. In ordinary metallurgical installations or pyro-metallurgical processes ore concentrates are first of all shifted from place to place and dust is created. In other cases the ores or concentrates are crushed or dried or roasted, and in each one of these operations dust and fumes, are created. All suspended particles and fumes originating from the abovementioned conditions not only, in many instances, provide health hazards which disrupt production and endanger the personnel in the plant, but also are the source of economic losses through extra maintenance and wear to moving machinery due to abrasion and corrosion. Further losses are incurred by the waste of fine, very often valuable, components. The Party who once said that in several metallurgical plants the bulk of the profits was discharged into the atmosphere by the stack wasn&apos;t stretching the truth. There is a place in Africa where a gold producing concern used to lose over 5 oz. of fine gold per day out of a total production of 3000 oz. per month. In another instance, copper concentrate going as high as 80% Cu was discharged into the atmosphere at the rate of 4 tons per day. Realising the value of removing and collecting solids and fumes created in metallurgical plants, a thorough study was made of the available plant and equipment offered on the market. Some units were found very efficient, but very expensive; some others also very efficient but with temperature and maintenance limitations. Some others were also efficient but prohibitive from the point of running costs.

Jan 1, 1963

By S. Kemball-Cook, D. Heinze, A. Kornberg, M. Harmon

Climate change presents risks to mine infrastructure, operations and long-term obligations to protect environmental conditions. Management strategies and design of infrastructure and containment systems based on past or current climate conditions may be inadequate in the future if changes in climate are not taken into account in planning for mine operations, decommissioning, closure and post-closure. Data-driven assessments using climate model projections of the future can be used to quantify: • Changes in site-specific risks to mine structures, access, containment facilities, operations and staff safety. • Changes in vegetation and water balance/chemistry, and potential for acid mine drainage, erosion and design failure. Climate- and weather-related risks are a serious threat to the mining industry, but risk transfer solutions can alleviate these vulnerabilities. Mining companies can protect themselves through bespoke solutions which provide operators, investors and buyers with protection against weather-related disruptions to operations by hedging to smooth revenues. Using risk transfer structures to mitigate climate risks is a mechanism that is available to all types of mining operations world-wide. In this paper, we describe methods for evaluating site-specific climate change impacts and developing risk transfer solutions to mitigate their effects. EVALUATING MINE RISKS FROM CLIMATE CHANGE The best available tool for understanding potential future changes to the earth’s climate system are global climate models (GCMs) (IPCC, 2014). GCMs are computer simulations of the climate system which include the atmosphere, ocean, land surface, cryosphere and biosphere. Because of the computational demands of running a global model for many decades, the spatial resolutions of the GCM grids are necessarily coarse, typically on the order of 100-300 km. GCM data are therefore suitable for evaluating future physical risks on global or continental scales, but may not be appropriate for evaluating location-specific risks (e.g. Kotamarthi et al., 2016). For site-specific, quantitative physical risk assessment, GCM projection data are “downscaled” or adapted for use at the local level using statistical techniques together with observed data (statistical downscaling) or using high-resolution regional climate models (RCMs) that focus on a single geographic region and have spatial resolutions of 10-50 km (dynamical downscaling). Downscaling bridges the gap between the GCM scale and regional/local scale. Through the downscaling process, new information is added to the GCM output from observations and/or RCMs, resulting in patterns for weather variables that are more consistent with local terrain and weather influences. Downscaled climate change projections for the mine site are selected considering mine location and geography, the availability of local observations, and the performance of the downscaling method and global/regional models used to develop the projection in simulating the variable under study. Where possible, we use projections from an ensemble of global and regional model runs spanning a range of future greenhouse gas (GHG) scenarios.

Jan 1, 2019

By R. Eggert

Since the discovery of high temperature superconductors has led to extensive interest and conjecture regarding their potential to change contemporary technology, discussions of the impact on the metallurgical and other industries have naturally arisen. In this paper we present data showing the possible impact of these developments on the materials related to the fabrication and proliferation of devices based on superconductor materials. We observe that the largest volume of effort with LTSC will be for the demonstration of high power devices while applied research relating to processing will remain the principal HTSC activity. We also note that in the next decade copper consumption in the US is expected to remain essentially at the current level while the need for Nb and Sn,; two major components of LTSC, will grow at a faster rate both in the US and the world. Although the use of superconductor-based magnetic separation is anticipated to increase, no major impact of these components on the metallurgical industries is expected.

Jan 1, 1990

By Andrea Brickey, RICHARD AMOAKO, Akshay Chowdu

Production forecasts derived from the medium-term schedule represent the best path forward for underground mining operations to achieve corporate and operational goals. Unforeseen circumstances, such as equipment breakdown, can cause significant deviations from established production forecasts in the short term. Such situations require timely mitigation of the effects of the deviations in order to stay as close as possible to the medium-term plan. This paper presents a deterministic mathematical formulation for short-term scheduling that addresses this challenge. The formulation is able to make up for deficits resulting from operational disruptions by minimizing deviations from medium-term production goals and medium-term activity start dates. We demonstrate this in a case study by simulating disruptions involving reduction in ore extraction capacity and development advance rate. Aside from the rescheduling capability, this formulation can help mine planners conduct analysis on the impact of various disruptions on the schedule. The formulation can help identify those conditions or disruptions that are more detrimental to the achievement of forecast targets, or otherwise. The formulation also affords mine planners the flexibility to prioritize one production goal over the other, depending on operational conditions at the mine. In the future, we seek to incorporate in the formulation respirable dust, diesel particulate matter and ventilation constraints in order to create more realistic schedules, and enhance miners’ health and safety.

Jan 1, 2024

By Diler Katircioglu-Bayel, Oner Yusuf Toraman, Safak Gokhan Ozkan

Calcite has found many applications in the industrialized countries recently, such as filler for paper, paint, plastics, rubber and food. Furthermore, calcite separates many toxic elements, and thus it is used in water treatment, and it is also used to raise the pH of drinking water. Due to its special features, calcite has been characterized as a useful coating and reinforcement filler for applications. Finer particle size of calcite has been determined to provide improved features for all of these applications. For submicron particle production, wet milling is selected over dry milling to produce finer particle sizes. Wet stirred media mills are also used in many industries. For that reason, this study investigated the effects of operating parameters on particle size reduction in the wet grinding process. The knowledge acquired by this study will allow the finding of optimal operating parameters for a stirred media mill to optimize specific energy consumption and reduce particle size.